Hydrometallurgical process for the recovery of precious metal values from precious metal ores with thiosulfate lixiviant

ABSTRACT

A hydrometallurgical process for the recovery of precious metal values from refractory precious metal ore materials containing preg-robbing carbon by leaching with a thiosulfate lixiviant, the process comprising 
     a. providing a body of particles and/or particulates of an ore material having precious metal values and preg-robbing carbonaceous components; 
     b. contacting the body of particles and/or particulates with a thiosulfate lixiviant solution at conditions conducive to the formation of stable precious metal thiosulfate complexes; 
     c. recovering the thiosulfate lixiviant from the body of particles and/or particulates after a period of contact which is sufficient for the lixiviant solution to become pregnant with precious metal values extracted from the ore material; and 
     d. recovering the precious metal values from the lixiviant solution. 
     In a preferred embodiment, the refractory ore materials are low grade precious metal ore materials.

This is a continuation-in-part of co-pending application Ser. No.07/861,563, filed on Apr. 1, 1992.

BACKGROUND OF THE INVENTION

The present invention relates to the hydrometallurgical recovery ofprecious metal values from refractory precious metal ores containingpreg-robbing carbonaceous material.

Conventionally, precious metals have been extracted from ore materialsby lixiviation or leaching with cyanide-containing solutions. It hasbeen found however that some gold ores do not respond well toconventional lixiviation because of the presence of impurities thatinterfere with the leaching process. These ores are termed "refractory".

One common cause of the refractoriness of gold ores is organiccarbonaceous matter that is associated with some deposits. Thiscarbonaceous matter is believed to adsorb solubilized gold complexesfrom lixiviant solutions back into the ore. The adsorbed gold is notrecovered, and remains with the ore material and is eventually carriedoff with the tailings, leading to poor gold recovery. This can be a veryserious problem, as a small amount of carbonaceous matter can adsorbessentially all of the solubilized gold in an entire cyanide lixiviationcircuit. This is sometimes referred to as poisoning the circuit. Inother cases, the carbonaceous matter is believed to coat the gold, andthereby prevent the lixiviant solution from gaining access to it. Inother words, carbon can "rob" precious metal values, and gold inparticular, from the lixiviant solution that is "pregnant" therewith.This characteristic is referred to as "preg-robbing."

It is believed that the carbonaceous content that participates inpreg-robbing comprises an activated carbon-type carbon material,long-chain hydrocarbons and organic acids, such as humic acid. SeeSibrell, P. L. et al., Spectroscopic analysis of Passivation Reactionsfor Carbonaceous Matter from Carlin Trend Ores, GOLD 90 PROCESSMINERALOGY X, pp. 355-363 (1990). Adsorption of the gold lixiviantcomplex by carbonaceous material is very complicated, due to three majorfactors. First, the precise chemical and physical nature of thecarbonaceous matter is difficult to define, and varies from one ore bodyto the next. Second, the mechanism by which the carbonaceous materialadsorbs gold is still being investigated. Third, although it has beenknown for some time that preg-robbing carbonaceous material can bepassivated, or treated so as not to adsorb gold, the mechanism by whichthis occurs is not fully understood.

Many procedures have been investigated in an effort to passivate ordeactivate the preg-robbing potential of carbonaceous ores, but nonehave been entirely satisfactory when applied to low grade refractoryores. The procedures heretofore tried include roasting, kerosenepretreatment, flotation, aqueous chlorination, chemical oxidation andbiological deactivation. The preferred approach for the recovery ofprevious metal values from preg-robbing carbonaceous ore materials hasbeen to deactivate or remove the preg-robbing components in the orematerial using one of the aforereferenced pretreatment techniquesfollowed by lixiviation with cyanide-solution. Examples of such attemptscan be found in U.S. Pat. No. 5,127,942 to Brierley et al. whichdescribes the deactivation the preg-robbing carbonaceous component inrefractory ores using a specific microbial consortium, followed byrecovery of precious metal from the carbon-deactivated residue bycyanidation; U.S. Pat. No. 4,801,329 to Clough et al. which describesthe use of a chemical oxidation pretreatment to enable a precious metalto be extracted from carbonaceous ores preferably by cyanidation. Thedeactivation of the preg-robbing carbonaceous components with chemicalagents, such as taught by Clough, and with biological/biochemicalagents, such as taught by Brierley, introduces additional expense andcomplexity to the processing of refractory ores materials.

Heretofore thiosulfate lixiviant has been suggested for recoveringprecious metals from difficult to treat ores. U.S. Pat. No. 4,654,078 toPerez et al. describes the use of copper-ammonium thiosulfate to recoverprecious metals from difficult-to-treat ores, especially thosecontaining manganese and/or copper. The presence of copper and/ormanganese contraindicates the use of cyanide solution leaching becausesuch materials increase cyanide consumption. U.S. Pat. Nos. 4,369,061and 4,269,622 to Kerley describe lixiviating with an ammoniumthiosulfate leach solution containing copper to recover precious metalsfrom difficult-to-treat ores, particularly those containing copper,arsenic, antimony, selenium, tellurium and/or manganese, and mostparticularly those containing manganese. U.S. Pat. No. 4,070,182 toGenik-Sas-Berezlosky et al. describes the recovery of gold fromcopper-bearing sulfidic material containing gold using a secondary leachwith ammonium thiosulfate.

Nothing in the prior art has suggested that excellent precious metalrecovery yields could be achieved from preg-robbing carbonaceous oresmaterial, including low grade materials, without a pretreatment step todeactivate or remove the preg-robbing components in the ore material byusing thiosulfate lixiviation under controlled conditions.

Despite the growing world-wide interest in recovering precious metalsfrom carbonaceous ores, and substantial work which has been done todevelop a viable technology for doing so, a fully satisfactory processfor metal recovery from most carbonaceous ore materials has yet to beprovided.

Therefore, it is an object of the invention to provide a process forrecovering at least one precious metal from preg-robbing carbonaceousore, without the necessity of first subjecting the preg-robbingcarbonaceous ore to a pretreatment step to deactivate or remove thepreg-robbing component of the ore.

A further object of the present invention is to permit the recovery ofprecious metals values from low grade precious metal refractory orematerial, including material that has heretofore been considered waste.

Still further, the present invention has as a goal the recovery ofprecious metal values from refractory precious metal ore material,particularly such ore materials with low grade precious metal content,with improved economic and energy efficiency.

SUMMARY OF THE INVENTION

The present invention relates to a hydrometallurgical process for therecovery of precious metal values from refractory precious metal orematerials containing preg-robbing carbonaceous material comprising:

a. providing a body of particles and/or particulates of an ore materialhaving precious metal values and a preg-robbing carbonaceous materialcontent;

b. contacting the body of particles with a thiosulfate lixiviantsolution at conditions conducive to the formation of stable preciousmetal thiosulfate complexes;

c. recovering the thiosulfate lixiviant solution from the body of orematerial particles and/or particulates after lixiviant solution ispregnant with precious metal values extracted from the ore material;

d. recovering the precious metal values from the lixiviant solution.

The invention can be practiced on a batch or continuous basis. Thecontacting step can take place on a pad, with the ore material to betreated situated in a heap; or in a vat, tank or the like. The primarycriterion for the contacting step is that the thiosulfate lixiviantsolution achieve intimate contact with the precious metal containing orematerial.

According to the invention, a refractory precious metal ore materialcontaining preg-robbing carbonaceous material is reduced by means ofcrushing, grinding or like processing to a particle size that isadvantageous for metallurgical liberation. The invention can be put intopractice in several different processing schemes and the particle sizeselected depends on which processing scheme is selected.

In one embodiment of the invention precious metal values are extractedfrom preg-robbing carbonaceous ore material in a heap. The term "heap"is used to describe a static body of ore material. It applies to a massof particles and/or particulates supported only at its base, such as aheap, etc., and also, if desired, to a mass of particles and/orparticulates supported on its sides in which the ore material remainsstatic such as being held in a confining vessel such as a column, vat,tank and the like--a form that is particularly advantageous forrecirculating a lixiviant solution.

When ore material is processed in a heap it is preferably in the form ofsubdivided particles and/or particulates with 90% by weight being lessthan 2 inches in size, and preferably 70-80% by weight being less than0.5 inch in size. By the term "particles" it is meant the individualparticles found in ore material, such as run-of-the-mine ore; further,it is meant, ore particles formed after crushing. By the term"particulates" it is meant a body or shape that is built up fromindividual particles properly agglomerated. Since the process of thepresent invention is particularly amenable to low grade ores and towaste, particles need not be milled or ground, thereby reducing thecapital and operating costs of the process of the present invention. Ifdesired, particulates can be formed or made. One method for formingparticulates is disclosed in U.S. Pat. No. 4,765,827 to Clough. Otherconventional methods for forming particulates include extruding,pilling, tableting and the like.

To facilitate the recovery of precious metal values from ores that havea sulfidic sulfur content that also renders them refractory, any suchsulfide content in the ore is preferably at least partially oxidized.Refractoriness in sulfidic ores is believed to be caused by very finegrain gold being encapsulated in sulfide minerals, such as pyrite,arsenopyrite or arsenian pyrite. In some cases, the gold occurs assubstitutional impurity atoms in the sulfide mineral crystal lattice.The sulfides must therefore be completely, or at least partially,oxidized to allow lixiviant solution access to the gold. It is desirablethat the sulfide content of such ores be decreased by about 40% or more.Suitable sulfide oxidation processes for refractory sulfidic preciousmetal ores are:

autoclaving,

chlorination,

nitric acid oxidation,

microbial oxidation (also known as biooxidation), or

roasting.

If sulfidic sulfur content of an ore material is treated by microbial ornitric acid oxidation, the ore material will be left with an acidic pH.In such cases, it is desirable to raise the pH of the ore to at leastabout 9 to enable the efficient recovery of precious metal values fromthe ore using thiosulfate lixiviation. This can be done by washing theore material and/or treating it with an aqueous solution having a basicpH. Sodium carbonate, dilute ammonium hydroxide, lime and caustic aresuitable bases.

Thiosulfate lixiviation of a static heap of ore material comprisespassing thiosulfate lixiviant in solution through the heap underconditions selected to cause the thiosulfate to extract precious metalvalues from the ore material. After passing through the heap, thethiosulfate lixiviant becomes pregnant with extracted precious metalsvalues. The pregnant lixiviant solution is recovered at the bottom ofthe heap and recirculated, either continuously or intermittently.Precious metal values are recovered from the lixiviant solution,preferably by means of precipitation. The recovery of precious metalsfrom the lixiviant solution can be carried out either periodically orcontinuously. After the precious metal values are recovered from thelixiviant solution, the regenerated solution is recirculated to thestatic heap.

It has been found that when the lixiviation of carbonaceous preg-robbingore material is conducted in accordance with the teachings of thepresent invention with respect to controlled thiosulfate concentration,lixiviant solution pH, oxidation/reduction conditions and ammoniaconcentration, as more fully described hereinafter, high precious metalrecovery yields can be achieved even from low grade preg-robbingcarbonaceous ore materials without pretreatment of the ore to deactivateor remove its preg-robbing component.

In an alternate embodiment of the invention, preg-robbing carbonaceousore materials containing precious metal values are finely ground beforebeing subjected to extraction of precious metal values using thiosulfatelixiviant. As described above with respect to the "static heap" process,it has been found that when thiosulfate lixiviant is used undercontrolled conditions, very good precious metal recovery yields can beachieved from ore materials containing preg-robbing carbonaceousmaterial without a pre-treatment step to deactivate or remove thepreg-robbing component.

Finely grinding ore material prior to subjecting it to thiosulfatelixiviation increases the surface area of the ore exposed to thelixiviating solution and achieves comparatively higher precious metalrecovery from a given ore material with less thiosulfate lixiviantcontact time. While higher precious metal recovery yield and shortenedextraction time are obvious benefits of this approach, finely grindingan ore material imposes additional capital burdens. It has historicallybeen justified only on higher grade ore materials. The selection betweenprocessing technology is generally made based on laboratory analysis ofthe refractory ore material.

Even though finely grinding refractory ore does tend to liberateprecious metals which are occluded in sulfide components within oreshaving high sulfidic sulfur contents, processing to reduce the sulfidecontent of such ores is often necessary just as aforedescribed withrespect to static heap processing. In such cases, it is within the scopeof the present invention to subject the finely ground ore material to apretreatment step to at least partially oxidize the sulfide sulfur inthe ore material. Oxidation pretreatments for sulfide sulfur in finelyground ore material include microbial oxidation, nitric acid oxidation,chemical oxidation and autoclaving. None of these sulfur pretreatmentsdeactivate or remove pre-robbing carbonaceous components in an orematerial. As aforedescribed, if the sulfidic sulfur content of an orematerial is treated by microbial or nitric acid oxidation, the finelyground ore material will be left with an acidic pH. In such cases, it isalso desirable to raise the pH of the finely ground ore material to atleast about 9 prior to contacting the ore material with thiosulfatelixiviant solution in the precious metal extraction step.

The finely ground preg-robbing ore, pretreated as aforedescribed, ifnecessary, to reduce refractory sulfide content, is slurried withthiosulfate lixiviant solution, and is preferably leached using a seriesof stirred contacting tanks through which the finely ground ore materialand the thiosulfate lixiviant solution flow countercurrently.Alternatively, precious metal values can be extracted from the finelyground preg-robbing ore material using thiosulfate lixiviant using batchprocessing. In either case, precious metal values are recovered from thelixiviant solution after the contacting step, preferably by means ofprecipitation.

BRIEF DESCRIPTION OF THE DRAWINGS

FIG. 1 is a schematic diagram of a precious metal value(s) lixiviationand recovery process for static heaps of ore material in accordance withthe present invention;

FIG. 2 is block diagram of the major steps in a precious metal value(s)lixiviation and recovery process for finely ground ore material inaccordance with the present invention;

FIG. 3 is a graph that plots the cumulative percent gold extracted fromthe ore sample versus column leaching duration in days for columns 1 and2 of Example 2;

FIG. 4 is a graph that plots the cumulative percent gold extracted fromthe ore sample versus column leaching duration in days for columns 3 and4 of Example 2;

FIG. 5 is a graph that plots the cumulative percent gold extracted fromthe ore sample versus column leaching duration in days at various cupricion concentrations in the lixiviant solution;

FIG. 6 is a graph that plots the cumulative percent gold extracted fromthe ore sample versus column leaching duration in days at various columnand lixiviant solution temperatures;

FIG. 7 is a graph that plots the cumulative percent gold extracted fromthe ore sample versus column leaching duration in days;

FIG. 8 is a graph that plots the cumulative percent gold recovered fromthe lixiviant solution versus elapsed process time in days forrecovering gold using zinc and copper cementation;

FIG. 9 is a graph that plots cumulative percent gold recovered from thelixiviant solution versus elapsed time in minutes using zinc cementationunder deaerated or atmospheric conditions; and,

FIG. 10 is a graph that plots cumulative percent gold recovered from thelixiviant solution versus elapsed time in minutes using coppercementation under deaerated or atmospheric conditions.

DETAILED DESCRIPTION OF THE INVENTION

As previously stated, the present invention is directed to ahydrometallurgical process for the recovery of precious metal valuesfrom a refractory precious metal ore materials containing preg-robbingcarbonaceous components comprising:

a. providing a body of particles and/or particulates of an ore materialhaving precious metal values and preg-robbing carbonaceous components;

b. contacting the body of particles and/or particulates with athiosulfate lixiviant solution at conditions conducive to the formationof stable precious metal thiosulfate complexes;

c. recovering the thiosulfate lixiviant from the body of particlesand/or particulates after a period of contact which is sufficient forthe lixiviant solution to become pregnant with precious metal valuesextracted from the ore material; and

d. recovering the precious metal values from the lixiviant solution.

The terms "ore" or "ore material" as used herein include not only oreper se, but also concentrates, tailings, spoil or waste in which asufficient level of precious metal value(s) exists to justify therecovery of those values. The present invention is particularlydesirable for use with low-grade ores and/or with materials consideredas waste.

Suitable candidate precious metal ores for the practice of the presentinvention are

1. mixed carbonaceous and sulfidic ores, such as carbonaceous-sulfidicores,

2. carbonaceous ores;

3. sulfidic ores, e.g., pyritic, arsenopyritic, or arsenian pyrite ores,in which the precious metal, e.g., gold, is associated with the sulfideand

4. mixtures of the foregoing in which preg-robbing carbonaceous materialis present.

Refractory carbonaceous-sulfidic and refractory carbonaceous oxide oreshaving a preg-robbing carbonaceous material content are candidate oresthat are amenable in an unexpected manner to the treatment according tothe present invention. The present invention is especially suitable forores that have a preg-robbing carbonaceous material content and enablesthe efficient recovery of gold from many low grade refractory ores fromwhich no gold or only small amounts of gold can be extracted withcyanide, even in laboratory bottle tests. Heretofore pretreatment todeactivate the preg-robbing carbonaceous content in such ores has beennecessary.

Specific ores that may be advantageously treated in accordance with thepresent invention are carbonaceous or carbonaceous-sulfidic or sulfidicores; for example, ores from the regions around Carlin, Nev.

A common characteristic of the deposits in the Carlin trend is thatsub-micron size gold is disseminated in a quartz or quartz/calcitematrix. Unoxidized ore zones contain organic carbonaceous matter andsulfidic minerals. These gold ores are refractory principally because ofthe carbonaceous matter contained in the ore. While sulfide minerals mayprevent access of a cyanide lixiviant solution to some of the gold, thecarbonaceous matter could poison an entire cyanide lixiviation circuit.To the extent that refractory characteristics of these ores derive fromtheir sulfidic content it can be satisfactorily handled with the sulfurpretreatments aforedescribed. To the extent, however, that refractorycharacteristics in these ore materials are occasioned by preg-robbingcarbonaceous components in the ore material, they are not costeffectively overcome in low grade ores by known carbon pretreatments,but are overcome by thiosulfate lixiviation in accordance with thepresent invention.

The thiosulfate lixiviation of the present invention comprisescontacting a body of particles and/or particulates of refractoryprecious metal ore material which contains preg-robbing carbonaceousmaterial with a solution of thiosulfate lixiviant under conditionproviding for intimate contact between the two.

When low grade ore material are processed in accordance with theinvention, the body of particles and/or particulates preferablycomprises a heap of agglomerated particles and particulates, formed asaforedescribed, and the contacting step preferably comprises passing thethiosulfate lixiviant solution through the heap by applying it to thetop of the heap at a controlled flow rate under conditions which causethe solution to flow through the heap and intimately wet theagglomerated ore particles. The thiosulfate lixiviant solution isrecovered at the bottom of the heap.

When higher grade ore material, containing relatively greater amounts ofprecious metal are processed in accordance with the invention, the bodyof particles and/or particulates of refractory precious metal orematerial may preferably comprise finely ground particles with a highpercentage having a grain size in the range of 200 mesh. In such casesthe contacting step preferably comprises forming a slurry of thiosulfatelixiviant solution and the finely ground ore in a stirred vessel.

The thiosulfate lixiviation of the invention can derive the necessarythiosulfate ion from a variety of sources, such as ammonium thiosulfateor sodium thiosulfate or a mixture of both, the conditions under whichlixiviation in accordance with the present invention occurs needs to becarefully controlled as to optimize thiosulfate stability, preciousmetal value extraction and complexing/solvation, and reagent management.In a preferred form, the lixiviant system has

1. an ammonium thiosulfate or sodium thiosulfate (or mixture of both)concentration of at least about 0.05M (corresponding to about 7.5 gramsof ammonium thiosulfate per liter of lixiviant solution) and preferablyfrom about 0.1M to about 0.2M (corresponding about 15 to about 30 gramsof ammonium thiosulfate per liter of lixiviant solution),

2. a pH of at least about 9 (and preferably about 9.2 to about 10)

3. an oxidizing agent, preferably cupric tetrammine ions Cu(NH₃)₄ ²⁺, insufficient concentration to catalyze the oxidation reaction, such asless than 0.001M (less than about 60 parts per million parts oflixiviant solution, and preferably from about 20 to about 30 parts) and

4. an ammonia concentration sufficient to stabilize the thiosulfatecomplex and the cupric tetrammine, such as at least about 0.05M andpreferably at least about 0.1M.

The overall stoichiometry for the dissolution of gold in aqueousthiosulfate solutions in the presence of oxygen is shown in Equation 1,as follows:

(Equation 1)

    2 Au+1/2O.sub.2 +4 S.sub.2 O.sub.3.sup.2- +H.sub.2 O ⃡2 Au(S.sub.2 O.sub.3).sub.2.sup.3- +2 OH.sup.-.

Cupric ion is believed to have a strong catalytic effect on the rate ofoxidation since the addition of cupric ion to a thiosulfate solutionresults in the formation of cupric thiosulfate, Cu(S₂ O₃)₃ ⁴⁻ or Cu(S₂O₃)₂ ²⁻ such that, in the presence of oxygen (such as air in the case ofheap leaching) the copper remains in an oxidation state as cupric ions.The presence of ammonium ions helps to stabilize the cupric oxidationstate as cuptic tetrammine complex ion--Cu(NH₃)₄ ²⁺. Not only does thepresence of ammonia facilitate the formation and stabilization of cuprictetrammine ions, but it aids in neutralizing the ore material andkeeping it alkaline. The role of cuptic tetrammine as an oxidant duringthe dissolution of gold is shown in Equation 2, as follows:

(Equation 2)

    Au+5 S.sub.2 O.sub.3.sup.2- +Cu (NH.sub.3).sub.4.sup.2+ ⃡Au(S.sub.2 O.sub.3).sub.2.sup.3- +4NH.sub.3 +Cu(S.sub.2 O.sub.3).sub.3.sup.5-

Equation (2) also depicts the cupric/cuprous equilibrium that exists inammoniacal thiosulfate solutions. However, under oxidizing conditions,oxidative degradation of thiosulfate to tetrathionate occurs and theoxidation reaction is promoted by cuptic ion which is described inEquation 3, as follows:

(Equation 3)

    2(NH.sub.4).sub.2 S.sub.2 O.sub.3 +1/2O.sub.2 +H.sub.2 O⃡(NH.sub.4).sub.2 S.sub.4 O.sub.6 +2NH.sub.4 OH.

Thus, the amount of cupric ion addition or the concentration of cupricion is an important factor in thiosulfate stability and reagentmanagement in the heap lixiviation of the present invention.

The pH value for the lixiviant solution (at least about 9) is alsoimportant in keeping the gold thiosulfate complex ion stable, i.e.,within the stabilized region as indicated in the corresponding Eh-pHdiagram for the lixiviant system.

After the lixiviant solution becomes pregnant with precious metal valuesby contacting the body of refractory precious metal ore material asaforedescribed, the precious metal values may be recovered from thelixiviant solution in a variety of ways, including preferably byprecipitation with:

a. copper (such as, metallic copper powder or a copper precipitate fromcementation),

b. zinc (such as, metallic zinc powder) or

c. aluminum (such as, metallic aluminum powder) or

d. soluble sulfides. In those instances where zinc, aluminum or solublesulfides are used as a precipitation agent, it may be desirable to addadditional copper ions (such as in the form of copper sulfate) in orderto maintain the desired level of cupric ion because zinc, aluminum andsoluble sulfide will also remove copper from the lixiviant solution.

It has long been recognized that the reduction of metals from solutionis a result of charge-transfer reactions. In this regard, the reactionfor gold recovery by zinc cementation can be presented in terms of therespective half cell reactions: cathodic reduction of gold thiosulfateanion as described in Equation 4, as follows:

(Equation 4)

    Au(S.sub.2 O.sub.3).sub.2.sup.3- +e.sup.- ⃡Au°+2 S.sub.2 O.sub.3.sup.2-

which is coupled to the anodic dissolution of zinc forming zincthiosulfate complex anion, or zinc ammonioum complex ion:

(Equation 5)

    Zn°+2 S.sub.2 O.sub.3.sup.2- ⃡Zn(S.sub.2 O.sub.3).sub.2.sup.2- +2e.sup.-, or

    Zn+4NH.sub.3 =Zn(NH.sub.3).sub.4.sup.2+ +2e.sup.-

The overall reaction for gold precipitation by zinc cementation in thethiosulfate solution is as described in Equation 6, as follows:

(Equation 6)

    Zn°+2 Au(S.sub.2 O.sub.3).sub.2.sup.3- ⃡2 Au°+Zn(S.sub.2 O.sub.3).sub.2.sup.2- +2 S.sub.2 O.sub.3.sup.2-, or

    Zn°+2 Au(S.sub.2 O.sub.3).sub.2.sup.3- +4NH.sub.3 =2 Au°+Zn(NH.sub.3).sub.4.sup.2+ +4S.sub.2 O.sub.3.sup.2-

Since cuptic and cuprous ions exist in the thiosulfate leachingsolution, the reduction reaction of cupric ion to cuprous ions by zincwill occur and the cuprous ion is further reduced to metallic copper.The cathodic reactions of Cu²⁺ /Cu⁺ and Cu⁺ /Cu are indicated inEquation 7 and 8 as follows: (Equation 7)

    2 Cu(S.sub.2 O.sub.3).sub.2.sup.2- +2e.sup.- ⃡CU.sub.2 (S.sub.2 O.sub.3).sub.3.sup.4- +S.sub.2 O.sub.3.sup.2-

and

(Equation 8)

    Cu.sub.2 (S.sub.2 O.sub.3).sub.3.sup.4- +2e.sup.- ⃡Cu°+3 S.sub.2 O.sub.3.sup.2-.

The overall reaction for copper precipitation by zinc can be expressedEquations 9 and 10 as follows:

(Equation 9)

    2Cu(S.sub.2 O.sub.3).sub.2.sup.2- +Zn°+S.sub.2 O.sub.3.sup.2- ⃡Cu.sub.2 (S.sub.2 O.sub.3).sub.3.sup.4- +Zn(S.sub.2 O.sub.3).sub.2.sup.2-

(Equation 10)

    Cu.sub.2 (S.sub.2 O.sub.3).sub.3.sup.4- +Zn°⃡2 Cu°+Zn(S.sub.2 O.sub.3).sub.2.sup.2- +S.sub.2 O.sub.3.sup.2-

Thus, the reactions for zinc cementation in the thiosulfate solutioninclude the reduction of (1) copper thiosulfate ion to metallic copperand (2) gold thiosulfate ion to metallic gold.

In the Merrill-Crowe zinc cementation process (using a cyanidelixiviation system), deaeration of the pregnant solution prior tocementation is one of the most important factors for efficient goldrecovery. The presence of oxygen passivates the surface of the zinc dustand also causes re-dissolution of the gold precipitate resulting inexcessive consumption of zinc and incomplete recovery of gold. However,in the thiosulfate lixiviation system of the present invention,deaeration in the cementation reaction system appears not to be criticalprovided sufficient zinc is added.

The process of the present invention, as applied to a static heap ofpreg-robbing ore material, is depicted schematically in FIG. 1 forpurposes of illustration, and not of limitation. As aforesaid, athiosulfate lixiviant solution is passed throughout heap 1 of theaforesaid particles and/or particulates, the precious metal-pregnantlixiviant solution is recovered and passed via conduit 10 to a pregnantlixiviant solution storage reservoir 20. Initially, it may be desirableto use the reservoir 20 to make up the initial lixiviant solution. Inthat case, water and ammonium thiosulfate are mixed together with asource of cupric ions, such as copper sulfate, sufficient to obtain theappropriate level of cupric ions and a source of ammonia, NH₃, to obtainthe appropriate level of ammonia concentration.

During operation of the lixiviation and recovery process, additionalcupric ions, ammonia and/or ammonium thiosulfate can be added to thereservoir 20 to maintain these reagents at their desired concentrations.Recovered lixiviant solution is moved from reservoir 20 through conduit40 to pump 30 and then through conduits 50, 60 and 70 to heap 1 where itis distributed over heap 1.

Precious metal recovery is effected by drawing a slipstream throughprecious metal recovery system 80, which may utilize zinc cementation.Recovered lixiviant solution is drawn from conduit 50 through conduit 90to recovery system 80 where the precious metal content of the recoveredpregnant lixiviant solution is partially or completely recovered and thelean lixiviant solution returned to the main stream conduit 70 viareturn slipstream conduit 100. Optionally, additional ammoniumthiosulfate (and other reagents, such as a source of cupric ions) can beadded to the slipstream after processing by recovery system 80. Ifdesired, flow through main stream conduit 60 can be cut off so that allpregnant lixiviant solution flow is through precious metal recoverysystem 80, so that the entire lixiviant solution is subjected toprecious metal recovery with each pass through the lixiviation andrecovery process; or it can be decreased so that there is partial flowthrough recovery system 80, whereupon there will be partial recovery ofprecious metal values with each pass of lixiviant solution through theprocess. Alternatively, the entire flow of pregnant lixiviant solutioncan be through main stream conduit 60 and none through the slipstream(conduits 90 and 100 and recovery system 80) until it is decided torecover precious metal value(s) from the pregnant lixiviant solution.

The thiosulfate lixiviant is passed throughout the heap, recovered andrecirculated, either continuously or intermittently, throughout thelixiviation stage. The thiosulfate lixiviant is recycled at apredetermined rate (for example from about 0.002 to about 0.01 gallonper minute per square foot of heap top surface area and preferably at arate of about 0.005 gallon per minute per square foot of heap topsurface area). It may be dispersed by means known in the art for heapleaching processes, with drip irrigation being the preferred dispersalmethod. Spraying the lixiviation on the heap can also be advantageous,since spraying can increase the oxygen content of the lixiviantsolution.

With each pass of the lixiviant solution through the heap theconcentration of solubilized precious metal value(s) in the solutionincrementally increases. While the precious metal value(s) present inthe solution may be recovered at the termination of the heap leachingprocess, it is preferred, and more efficient, to recover the preciousmetal values either continuously or intermittently during heap leaching,since a greater amount of precious metal can be leached out of the orematerial when the lixiviant solution is not heavily loaded with preciousmetal values. Further, it is preferable to recover the precious metalvalues from a portion of the recycle stream of lixiviated solution,using the slipstream aforedescribed. Another approach to precious metalrecovery is to continue the thiosulfate lixiviant recirculation untilthe precious metal content of the lixiviant no longer increases witheach recycle and to then recover the precious metal values from thethiosulfate lixiviant.

The process of the invention, as applied to finely ground refractoryprecious metal ore material contain preg-robbing carbonaceous componentsis illustrated in FIG. 2.

Finely ground carbonaceous ore material 101 is initially slurried withwater 102, thiosulfate lixiviant 103, copper sulfate 104 and ammonia 105in slurry preparation unit 110. Each reagent is added in appropriatequantity to establish a lixiviant solution having thiosulfateconcentration, solution pH, oxidizing agent concentration and ammoniaconcentration in the ranges aforedescribed. If finely groundcarbonaceous ore material 101 contains sulfidic sulfur which renders orematerial 101 refractory, such sulfide content is optionally at leastpartially oxidized in sulfur pretreatment unit 120 before ground ore 101reports to slurry preparation unit 110. Sulfur pretreatment isaccomplished by conventional means, with preferred modes being microbialoxidation, nitric acid oxidation or autoclaving. The slurry produced inslurry preparation unit 106 has a solids content of between 30 and 60weight percent, and preferably a solid content between 40 and 50 weightpercent.

During the operation of the lixiviation and recovery process, strippedthiosulfate lixiviant 161 is pumped from precious metal recovery unit160 into slurry preparation unit 110, and provides a significant part ofthe water, thiosulfate lixiviant, cupric ion and ammonia required inground ore slurry 111. Consequently, in normal operation, only make-upquantities of water 102, thiosulfate lixiviant 103, copper sulfate 104and ammonia 105 are added to slurry preparation unit 110 to achieve thedesired levels thereof in the lixiviation circuit.

Ground ore slurry 111 is transferred to heat exchanger 130 where, ifnecessary, the temperature of slurry 111 is adjusted to between about20° C. and about 45° C. and preferably between about 25° C. and about35° C. Higher temperatures within the given range have been found toincrease the percentage of available gold extracted from mostpreg-robbing carbonaceous ore materials, but also to increase thethiosulfate losses during lixiviation.

Slurry 111 is next passed to a stirred tank reactor 140 where thethiosulfate lixiviant operates to extract gold, or other precious metalsvalues, from the ground carbonaceous ore material 101. This extractionoperation can be carried out in a single stage, or in a plurality ofstages, wherein the extracted ore from the first stage is advanced to asucceeding stage, but the thiosulfate lixiviant preferably flows in acounter-current path from the last extraction stage to the first. Thelixiviant circuit depicted in FIG. 2 comprises a single stage, however,the number of lixiviant stages can range from a minimum of one stage tofour or more stages.

Ground ore slurry 111 is held in stirred tank reactor 140 until thethiosulfate lixiviant in slurry 111 extracts the desired amount of gold,or other precious metal values, from carbonaceous ore material 101. Whenthe lixiviation circuit is operated in the preferred conditions of theinvention, the interval of time required for this extraction process isprincipally controlled by the composition of the ore material beinglixiviated, the grain size of the ore material and the number oflixiviation stages in the circuit. Generally, the more finely an orematerial has been ground, the shorter the required extraction interval.Ordinarily, the total extraction time in a lixiviation circuit accordingto the invention will be between 2 and 18 hours and preferably between 4and 8 hours.

After the appropriate contacting interval, lixiviated ore slurry 141 istransferred from stirred tank reactor 140 to separator 150. Separator150 is of conventional design wherein the separator overflow is thepregnant lixiviant 151 and the underflow leached residue 152. Residue152 is transferred to tailings.

Precious metal recovery is effected by processing pregnant thiosulfatelixiviant 151 in precious metal recovery unit 160. Precious metal valuesare separated from pregnant thiosulfate lixiviant 151, preferably bymeans of precipitation, such as by zinc cementation or coppercementation. The stripped lixiviant solution 161 is returned to thelixiviation circuit by pumping it to slurry preparation unit 110, whereit provides a substantial portion of the water and reagent requirementsneeded to convert ground carbonaceous ore material 101 into ground oreslurry 111. Additional thiosulfate lixiviant, copper sulfate to providecupric ions and/or ammonia are added to stripped thiosulfate solution161 as required to obtain the lixiviant solution pH and oxidizing agent,ammonia and lixiviant concentration previously described. Precious metalstream 162 is removed from precious metal recovery unit 160 and can befurther refined using standard techniques.

In the examples to follow various aspects of the present invention arefurther amplified and such amplifications are intended to beillustrations, but not limitations, of the invention disclosed herein.

EXAMPLE 1

The ore tested was a Gold Quarry carbonaceous/sulfidic ore ("GQ C/S"ore). It had been biooxidized in a sulfur biooxidation test heap forabout three months. Particle size during biooxidation ranged from 3inches to minus 30 mesh. The sample after biooxidation was mixed wellfor testing. Feed samples before biooxidation and after biooxidationwere assayed and the results were averaged and appear in Table 1.

                  TABLE 1                                                         ______________________________________                                        Chemical Analysis of Gold Quarry Carbonaceous/Sulfidic Ore,                   Before Biooxidation and After Biooxidation                                                          Sample from Test                                                    Sample Before                                                                           Heap (After                                                         Biooxidation                                                                            Biooxidation)                                           ______________________________________                                        Au, opt       0.064-0.078  0.068                                              Au* (CN), opt  0.003       0.001                                              Au PR value, opt                                                                             0.070      NA                                                  C (total), %  1.25        1.22                                                C (organic), %                                                                              1.20        1.20                                                S (total), %  2.36        1.80                                                S (sulfate), %                                                                              1.04        0.95                                                S (sulfide), %                                                                              1.32        0.85                                                Iron, %       1.58        NA                                                  ______________________________________                                         *Cyanide leachable gold.                                                 

Test heap biooxidation resulted in sulfide oxidation of 35%.

Results for Au(CN) indicate the gold in the sample that is leachable bycyanide. The ratio of Au(CN) to Au is 0.015 which indicates that only1.5% can be extracted by cyanidation--even after biooxidation.Therefore, the sample is very refractory.

The sample was submitted for semiquantitative X-ray diffraction analysiswhich indicated that the sample comprised 72% quartz, 10% alunite, 7%sericite, 4% kaolin, 3% barite and 3% iron oxides by weight.

The sample was crushed to minus 10 mesh. Laboratory columns were loadedwith 500 grams of this minus 10 mesh material, washed thoroughly withwater and conditioned with Na₂ CO₃ solution 10 grams per liter to adjustthe pH of the ore sample. Leach solution was continuously recycled tothe top of the column and dripping from top to bottom. Two column leachtests were conducted; one column for ammonium thiosulfate and one forsodium thiosulfate. Both column tests were performed at a pH of about10, using 0.085M thiosulfate solution, 0.01M of cupric ion and weightratio of liquid to solid of 2. Test results are shown in Table 2 asfollows.

                  TABLE 2                                                         ______________________________________                                        Thiosulfate Column Leach Tests on 10 Mesh Sample                                            Ammonium                                                        Reagent       Thiosulfate                                                                             Sodium Thiosulfate                                    ______________________________________                                        Pregnant: Au ppm                                                              Time:                                                                         24 hours      0.472     0.403                                                 48 hours      0.488     0.473                                                 72 hours      0.506     0.518                                                 Residue: Au, opt                                                                            0.024     0.030                                                 Calculated Head:                                                                            0.062     0.068                                                 Au, opt                                                                       Au Extraction at                                                              72 hours                                                                      (1)           64.71%    55.88%                                                (2)           61.29%    55.88%                                                ______________________________________                                         (1) Based on head and residue assay                                           (2) Based on calculated head                                             

Gold extraction based on residue assay after 72 hours was 64.7% byammonium thiosulfate leach and was 55.9% by sodium thiosulfate leach.

The column leach tests were conducted by recycling leach liquor withoutgold recovery. The long retention time with increasing goldconcentration in pregnant solution confirmed that gold thiosulfatecomplex was not adsorbed by refractory carbonaceous material in the oreunder these conditions.

EXAMPLE 2

Gold Quarry carbonaceous/sulfidic ore from the same source as Example 1and having the same composition as indicated in Example 1, after threemonths of biooxidation on the test heap as previously described, wasstage crushed with 100% passing 1/2" and then further biooxidized in alaboratory column for about 2 weeks. The sulfide oxidation of thissample was about 47%. Gold extraction using cyanide lixiviation was lessthan one percent (1%) indicating that the ore was very refractory.

Columns used in this study were 2 inches in diameter and 12 inches inheight. Each column was loaded with 500 grams of sample. Thebiooxidation sample was washed with water to remove most of thesolubilized iron and conditioned with Na₂ CO₃ solution 10 grams perliter for about 2 days. Ammonium thiosulfate solution was prepared asspecified by the testing. Solution pH was adjusted in the range of from9 to 10 with sodium carbonate. Leach solution was pumped at a flow rateof 2.0 ml/min from the solution reservoir to the top of the column,collected from the bottom of the column into the same reservoir andrecycled back to the leach column. No gold recovery unit was connectedto the leach system. Solution samples were taken from the leachreservoir on a timed interval and submitted for gold and copperanalysis. Thiosulfate concentration was measured as well as pH and Ehvalue of these solution samples. Thiosulfate concentration wasdetermined by a iodometric titration method in which a solution sampleis titrated with standard iodine solution at controlled pH. In somecases, solution was treated with formaldehyde to fix sulfite. A standardthiosulfate solution was prepared to calibrate the iodine solutionbefore titration.

Following leaching, residues from these column tests were sampled,pulverized and submitted for chemical analysis. Gold extractions werecalculated based on the residue and feed assay data.

Column tests were performed with various concentrations of ammoniumthiosulfate in the range of 8.7 to 16.9 gpl. Because no gold recoverysystem was included in these column leach tests, the ammoniumthiosulfate solution was periodically changed with freshly preparedsolution. This was done to maintain a strong driving force for goldleaching and to simulate the situation in which a gold recovery systemwould be included. The ammonium thiosulfate leach conditions and resultsfor the column tests (including the results for each period in whichfresh thiosulfate solution had been added) are presented in Table 3.

                  TABLE 3                                                         ______________________________________                                        Ammonium Thiosulfate Leach Column                                                        Column Column   Column   Column                                               1      2        3        4                                         ______________________________________                                        (NH.sub.4).sub.2 S.sub.2 O.sub.3, gpl                                                      8.44     16.87    14.8   14.8                                    Liquid/Solid Ratio                                                                         0.5/1    0.5/1    1/1    1/1                                     pH           ˜9.8                                                                             ˜9.8                                                                             ˜9.9                                                                           ˜9.3                              Total leach, Days                                                                          15.8     15.8     12.3   12.3                                    Period I                                                                      Leach Days   8.8      8.8      3.3    3.3                                     Au Extraction %                                                                            27.37    45.53    27.72  38.16                                   Period II                                                                     Leach Days   7.0      7.0      5.0    5.0                                     Au Extraction %                                                                            15.28*   12.21*   15.85  21.42                                   Period III                                                                    Leach Days   --       --       4.0    4.0                                     Au Extraction %                                                                            --       --       6.44*  8.07                                    Residue                                                                       Au, opt      0.039    0.027    0.034  0.022                                   Cu, ppm      150.0    135.0    163.46 158.66                                  S (total), % 1.79              1.70   1.78                                    S (sulfide), %                                                                             0.50     1.70     0.44   0.40                                                          0.51                                                    Au Extraction, %                                                                           42.65    60.29    50.00  67.65                                   ______________________________________                                         *Included gold in wash solution.                                         

Based on daily leach solution assay data, cumulative gold extractions ofthese columns are presented in FIGS. 3 and 4. The thiosulfate solutionwas replaced with fresh thiosulfate solution during the test. Withrespect to columns 1 and 2, the thiosulfate solution was replaced atabout nine days and then drained at about 16 days followed by a waterwash. With respect to columns 3 and 4, the thiosulfate solution wasreplaced at about three days, again at about eight days and then drainedat about 12 days followed by a water wash. As can be seen from FIGS. 3and 4 the gold concentration is affected by the thiosulfateconcentration and liquid-to-solid ratio. One important observation fromthe FIGS. 3 and 4 results is that the gold concentration did notdecrease after recycling the pregnant solution during the same leachperiod. This indicates that gold thiosulfate was not preg-robbed by theore's carbonaceous species under these conditions.

Among these tests, column 4 gave the highest gold extraction (67.65%)with residue gold of 0.022 opt (ounces per ton). The leaching solutionvolume used in columns 3 and 4 was twice the amount used in columns 1and 2, and one addition of fresh thiosulfate solution was used forcolumns 3 and 4 during the leach period. Column 2 gave gold extractionof 60% with residue gold of 0.027 opt.

EXAMPLE 3

Samples from the same ore as EXAMPLE 2 and having the same treatment asEXAMPLE 2 were column leached under the same conditions as EXAMPLE 2were column leached under the same conditions as EXAMPLE 2, except thatan ammonium thiosulfate lixiviant solution having 15 grams per liter ofammonium thiosulfate, an ammonia (NH₃) concentration of 1.7 grams perliter and a pH of approximately 9.2 was used. The cuptic ionconcentration was either 0.0005 Molar (31.8 milligrams per liter) or0.001 Molar (63.5 milligrams per liter.) The results are shown in FIG. 5in which the cumulative percentage of gold extraction is plotted againstthe duration of time the respective lixiviant solutions are recirculatedthrough the column. It was determined that a lower concentration ofcuptic ions (0.0005M) was more effective when used in column leaching.

EXAMPLE 4

Samples from the same ore as EXAMPLE 2 and having the same treatmentwere column leached under the same conditions as EXAMPLE 2, except thateach of three columns and lixiviant solutions were operated at 5° C.,22° C. and 45° C. to determine the effect of ambient temperature on thelixiviant of gold. The results are shown in FIG. 6 in which cumulativepercent of gold extraction is plotted against the duration of time thelixiviant solution was recirculated through a column at each of therespective temperatures. FIG. 6 demonstrates that thiosulfatelixiviation according to the present invention is very robust withrespect to ambient temperatures.

EXAMPLE 5

Samples were taken from the same sample ore as Example 2 and had thesame composition. Sulfide oxidation of this sample was 47%. Followingbiooxidation, the sample was washed with water and agglomerated with 5lb/ton of cement. Thiosulfate leaching was conducted at a concentrationof 0.2M (or about 30 gpl) ammonium thiosulfate, 0.1M of free ammonia and0.0003M of cupric ion. 45.5 kilograms of sample was leached in a 8 inchcolumn. Solution to solid sample ratio was 0.2:1 and the leach solutionflow rate was controlled at 0.005 gallon per minute per square foot(GPM/ft²). The pregnant solution was periodically replaced with freshsolution. After 24 days of leaching, the samples were taken out of thecolumn, mixed and submitted for chemical analysis. The average chemicalanalytical results of this leach residue are given in Table 5, asfollows:

                  TABLE 5                                                         ______________________________________                                               Au   S (total)  S (sulfate)                                                                             S (sulfide)                                         opt  %          %         %                                            ______________________________________                                        Leach    0.030  1.68       1.31    0.37                                       Residue                                                                       ______________________________________                                    

Gold extraction based on residue and leach analysis was 55.2% after 24days leaching. The gold extraction curve is given in FIG. 7.

EXAMPLE 6

Thiosulfate column leaching was tested on a partially biooxidized sampleof Gold Quarry Carbonaceous/Sulfidic ore obtained from anotherbiooxidation heap at a particle size of about 1 inch. The sample wascollected after 95 days of biooxidation with an average sulfide sulfurcontent of 0.43% as compared to 0.83% sulfide sulfur in the feed sample(before biooxidation). About 48% of the sulfide content was oxidizedafter biooxidation. The average chemical composition analytical data ofthe sample before and after biooxidation are given in Table 6.

                  TABLE 6                                                         ______________________________________                                        Chemical Analysis of Carbonaceous/Sulfide                                     Ore Before Biooxidation and After Biooxidation                                                      Sample After                                                      Sample Before                                                                             Biooxidation                                                      Oxidation   Test 1  Test 2                                          ______________________________________                                        Au, opt     0.073         0.074   0.078                                       Au (CN), opt                                                                              0             0       0                                           C (organic), %                                                                            1.00          1.02    0.97                                        S (total), %                                                                              1.74          1.80    1.48                                        S (sulfate), %                                                                            0.92          1.37    1.18                                        S (sulfide), %                                                                            0.83          0.43    0.50                                        ______________________________________                                    

Cyanide gold extraction Au(CN) was about zero indicating the highlyrefractory character of this sample.

A total amount of 91 kilograms or 200 lbs. of sample was used for eachtest. Each ore sample was washed with water and agglomerated with 5lb/ton cement before thiosulfate column leaching. Leach solution waspumped from a reservoir (with 10 liters of lixiviate) to the top of thecolumn at a flow rate of 0.005 gpm/ft². The effluent was analyzed andperiodically gold was removed from the solution by cementation process.Barren solution was recycled for further leaching.

Two tests are presented as an example. Thiosulfate leaching wasconducted at an initial concentration of 0.2M ammonium thiosulfate, 0.1Mof free ammonia and 0.003M of cupric ion.

In test 1, the gold in the pregnant solution was periodically removedfrom solution by cementation with zinc powder. After zinc cementation,the barren solution was returned to the reservoir in the leach circuit.The thiosulfate and copper concentrations were adjusted based on theanalytical data.

In test 2, the gold in the pregnant solution was periodically removedfrom solution by cementation with copper powder. Also, only half of thepregnant solution was split to gold recovery. After cementation, thebarren solution was combined with the pregnant solution and returned tothe column for recirculation.

Gold extractions for test 1 and test 2 are shown in FIG. 8. While itwould appear that the rate of extraction for test 2 is slower than thatfor of test 1, this is only an experimental artifact created by thedifferent cycle times involved with the cementation process. Theextractions were nearly identical after 14 days of leaching.

EXAMPLE 7

This example presents the test results for gold recovery by zinccementation from thiosulfate solution. The cementation of gold wastested from an actual column leach solution with Au 1.27 ppm, S₂ O₃ ²⁻13.82 grams per liter, Cu 28.9 ppm and pH˜9.3 at 23° C. The amount ofzinc powder addition was 0.23 grams per liter. Gold recovery as afunction of time was compared under deaeration conditions and alsoatmospheric conditions, i.e., open to the air. Results are given in FIG.9. In both cases, the gold cementation reaction was very fast andcomplete gold precipitation was achieved within 10 minutes.Precipitation of the copper behaves similarly to that of the gold.

EXAMPLE 8

Example 7 was repeated, except that gold recovery was performed by Cucementation under deaerated condition and open to the air with 0.13grams per liter of copper powder being used instead of zinc powder, athiosulfate concentration of 15 grams per liter and a temperature of 22°C. The results are illustrated in FIG. 10 which shows virtually nodifference between cementation under deaerated conditions and aeratedconditions.

While the present invention has been discussed with respect to sulfidicand mixed carbonaceous and sulfidic ore materials in which preg-robbingcarbon is present, it should be understood that the present invention isalso amenable to use with carbonaceous ore materials in whichpreg-robbing carbon is present.

While the exact reasons that cause the process of the present inventionto produce the herein-observed results are not fully known and could notbe predicted, the results themselves bespeak the achievements that havebeen obtained-based merely on the percent of gold extraction from theserefractory ores at improved economies and using a less complicatedapproach than prior technology can show.

It is also evident from the foregoing that various combinations andpermutations may we be practiced and advanced, but these are not to beunderstood as limiting the invention which has been defined in theclaims which follow.

We claim:
 1. A hydrometallurgical process for the recovery of preciousmetal values from a refractory precious metal ore material comprisinga.providing a static heap of particles and/or particulates comprising anore material having precious metal values and a preg-robbing carboncontent said heap having a pH of at least about 9; b. passing athiosulfate lixiviant solution through the static heap of particlesand/or particulates; c. recovering the thiosulfate lixiviant solutionpregnant with extracted precious metal values after it has passedthrough the static heap; d. recycling the recovered thiosulfatelixiviant solution to step b.; and e. at least periodically recoveringthe precious metal values from the thiosulfate lixiviant solution.
 2. Ahydrometallurgical process as defined by claim 1 wherein said preciousmetal values comprise gold.
 3. A hydrometallurgical process as definedby claim 1 wherein said ore material is selected from at least onemember of the group consisting of mixed sulfidic and oxide ores,carbonaceous ores and sulfidic ores.
 4. A hydrometallurgical process asdefined by claim 1 wherein 90% by weight of said particles and/orparticulates are less than two inches in size.
 5. A hydrometallurgicalprocess as defined in claim 1 wherein said thiosulfate lixiviantsolution has a pH of at least about
 9. 6. A hydrometallurgical processas defined by claim 1 wherein said thiosulfate lixiviant solutioncomprises an aqueous solution of ammonium thiosulfate or sodiumthiosulfate or a mixture of both.
 7. A hydrometallurgical process asdefined by claim 6 wherein said thiosulfate is present in aconcentration of at least .05M.
 8. A hydrometallurgical process asdefined by claim 1 wherein said lixiviant solution further comprisescupric ions.
 9. A hydrometallurgical process as defined by claim 8wherein said cupric ions are present in a concentration from about 20 toabout 30 parts per million parts of lixiviant solution.
 10. Ahydrometallurgical process as defined by claim 9 wherein said cupricions are present in the form of tetrammine cupric ions.
 11. Ahydrometallurgical process as defined by claim 1 wherein said lixiviantsolution further comprises free ammonia.
 12. A hydrometallurgicalprocess as defined by claim 11 wherein said ammonia is present in aconcentration of at least about 0.05M.
 13. A hydrometallurgical processas defined by claim 1 wherein said precious metal values are recoveredfrom the thiosulfate lixiviant solution by zinc cementation, coppercementation, aluminum cementation or soluble sulfide precipitation. 14.A hydrometallurgical process as defined by claim 1 wherein said preciousmetal values are recovered from the thiosulfate lixiviant solution bycementation.
 15. A hydrometallurgical process as defined by claim 10wherein said precious metal values are recovered from the thiosulfatelixiviant solution by either copper or zinc cementation.
 16. Ahydrometallurgical process as defined by claim 1 wherein precious metalvalues are recovered continuously from the thiosulfate lixiviantsolution.
 17. A hydrometallurgical process as defined in claim 1 whereinthe entire recovered thiosulfate lixiviant solution is subjected torecovery of precious metal values before recycling to step b.
 18. Ahydrometallurgical process as defined by claim 1 wherein a portion ofthe recovered thiosulfate lixiviant solution is subject to recovery ofprecious metal values before recycling to step b.
 19. Ahydrometallurgical process as defined by claim 1 wherein said orematerial comprises a sulfidic ore and said ore material has beenbiooxidized with a microbial agent to decrease the sulfidic sulfurcontent of said ore material.
 20. A hydrometallurgical process asdefined in claim 19 wherein at least about 40% of the sulfidic sulphurcontent of said sulfidic ore has been biooxidized with said microbialagent.
 21. A hydrometallurgical process as defined by claim 1 whereinsaid static heap has an upper surface defining a top of said heap andsaid thiosulfate lixiviant solution is recycled to step b. at a ratefrom about 0.002 to about 0.01 gallon per minute per square foot ofsurface area at the top of said static heap.
 22. A hydrometallurgicalprocess as defined by claim 1 wherein about 90% by weight of saidparticles and/or particulates are less than 2 inches in size, saidthiosulfate lixiviant solution comprises an aqueous solution of ammoniumthiosulfate or sodium thiosulfate or a mixture of both, and saidsolution has a thiosulfate concentration of at least 0.05M and a pH ofat least about
 9. 23. A hydrometallurgical process as defined by claim22 wherein said precious metal values comprise gold and said lixiviantsolution further comprises cupric ions in a concentration from about 20to about 30 parts per million and free ammonia in a concentration of atleast about 0.05M.
 24. A hydrometallurgical process as defined by claim23 wherein a portion of the recovered thiosulfate lixiviant solution issubjected to recovery of precious metal values before recycling to stepb.
 25. A hydrometallurgical process for the recovery of precious metalvalues from a static heap of particles and/or particulates of orematerial containing precious metal values and preg-robbing carbonaceousmaterials which have not been deactivated with chemical agents orbiological agents, comprising: adjusting the pH of the static heap ofore materials to at least about 9, extracting at least a portion of saidprecious metal values from the static heap of ore materials by passing athiosulfate lixiviant solution through said heap, recovering thethiosulfate lixiviant solution pregnant with extracted precious metalvalues after it has passed through said static heap, recirculating atleast a portion of the recovered thiosulfate lixiviant solution throughsaid static heap, and at least periodically recovering said preciousmetal values from said thiosulfate lixiviant solution.
 26. Ahydrometallurgical process as defined by claim 25 wherein saidthiosulfate lixiviant solution comprises an aqueous solution of ammoniumthiosulfate or sodium thiosulfate or a mixture of both having athiosulfate concentration of at least 0.05M and a pH of at least
 9. 27.A hydrometallurgical process as defined by claim 26 wherein saidprecious metal values comprise gold, said lixiviant solution furthercomprises cupric ions in a concentration from about 20 to about 30 partsper million and free ammonia in a concentration of at least about 0.05M.28. A hydrometallurgical process as defined by claim 26 wherein aportion of the thiosulfate lixiviant solution pregnant with extractedprecious metal values has at least a portion of said precious metalvalues recovered therefrom and is then recirculated through said staticheap.
 29. A hydrometallurgical process as defined by claim 26 whereinsaid static heap has an upper surface defining the top of said staticheap, about 90% by weight of said particles and/or particulates in saidstatic heap are less than 2 inches in size, and said thiosulfatelixiviant solution is passed through said static heap at a flow ratefrom about 0.002 to about 0.01 gallons per minute per square foot ofsurface area at the top of said static heap.
 30. A hydrometallurgicalprocess as defined by claim 25 wherein said ore material is selectedfrom at least one member of the group consisting of mixed sulfidic andoxide ores, carbonaceous ores and sulfidic ores.
 31. Ahydrometallurgical process as defined by claim 30 wherein said orematerial comprises a sulfidic ore having sulfidic sulphur content andsaid ore material has been biooxidized with a microbial agent todecrease the sulfidic sulphur content of said ore material.
 32. Ahydrometallurgical process for the recovery of precious metal valuesfrom a refractory precious metal ore material containing precious metalvalues and preg-robbing carbonaceous compounds comprising:a. providing abody of particles and/or particulates of the refractory precious metalore material; b. contacting the body of particles and/or particulateswith a thiosulfate lixiviant solution at conditions conducive to theformation of stable precious metal thiosulfate complexes; c. recoveringthe thiosulfate lixiviant from the body of particles and/or particulatesafter a period of contact which is sufficient for the lixiviant solutionto become pregnant with precious metal values extracted from the orematerial; and d. recovering the precious metal values from the lixiviantsolution.
 33. A hydrometallurgical process as defined by claim 32wherein said ore material is selected from at least one member of thegroup consisting of mixed sulfidic and oxide ores, carbonaceous ores andsulfidic ores.
 34. A hydrometallurgical process as defined by claim 33wherein said thiosulfate lixiviant solution comprises an aqueoussolution of ammonium thiosulfate or sodium thiosulfate or a mixture ofboth having a thiosulfate concentration of at least 0.05M and a pH of atleast
 9. 35. A hydrometallurgical process as defined by claim 34 whereinsaid precious metal values comprise gold, said lixiviant solutionfurther comprises cupric ions in a concentration from about 20 to about30 parts per million parts of lixiviant solution and free ammonia in aconcentration of at least about 0.05M.
 36. A hydrometallurgical processas defined by claim 35 wherein said cupric ions are present in the formof tetrammine cupric ions.
 37. A hydrometallurgical process as definedby claim 33 wherein said ore material comprises a sulfidic ore havingsulfidic sulphur content and said ore material has been biooxidized witha microbial agent to decrease the sulfidic sulphur content of said orematerial.
 38. A hydrometallurgical process as defined in claim 37wherein at least about 40% of the sulfidic sulfur content of saidsulfidic ore has been biooxidized with said microbial agent.
 39. Ahydrometallurgical process as defined by claim 32 wherein said preciousmetal values are recovered from the thiosulfate lixiviant solution byzinc cementation, copper cementation, aluminum cementation or solublesulfide precipitation.
 40. A hydrometallurgical process as defined byclaim 32 wherein said body of particles and/or particulates ofrefractory precious metal ore material comprises a static heap and saidstatic heap has an upper surface defining the top of said static heap,and said thiosulfate lixiviant solution is passed through said staticheap at a flow rate from about 0.002 to about 0.01 gallons per minuteper square foot of surface area at the top of said static heap.
 41. Ahydrometallurgical process as defined by claim 40 wherein 90% by weightsaid particles and/or particulates are less than two inches in size. 42.A hydrometallurgical process as defined by claim 41 wherein saidthiosulfate lixiviant solution comprises an aqueous solution of ammoniumthiosulfate or sodium thiosulfate or a mixture of both having athiosulfate concentration of at least 0.05M and a pH of at least
 9. 43.A hydrometallurgical process as defined by claim 42 wherein said orematerial is selected from at least one member of the group consisting ofmixed sulfidic and oxide ores, carbonaceous ores, and sulfidic ores. 44.A hydrometallurgical process as defined by claim 43 wherein said orematerial comprises a sulfidic ore having sulfidic sulfur content andsaid ore material has been biooxidized with a microbial agent todecrease the sulfidic sulfur content of said ore material.
 45. Ahydrometallurgical process as defined by claim 44 wherein said preciousmetal values comprise gold and said lixiviant solution further comprisescupric ions in a concentration from about 20 to about 30 parts permillion parts of lixiviant solution.
 46. A hydrometallurgical process asdefined by claim 44 wherein a portion of the thiosulfate lixiviantsolution6 pregnant with extracted precious metal values has at least aportion of said precious metal values recovered therefrom and is thenrecirculated through said static heap.